AU2017279746B2

AU2017279746B2 - Beneficiation of Lead Sulphide Bearing Material

Info

Publication number: AU2017279746B2
Application number: AU2017279746A
Authority: AU
Inventors: Chung Ho Lam; Andrew Robert Tong
Original assignee: Minicap Holdings Pty Ltd
Current assignee: Minicap Holdings Pty Ltd
Priority date: 2016-12-21
Filing date: 2017-12-21
Publication date: 2023-11-09
Anticipated expiration: 2037-12-21
Also published as: CN108315565A; AU2017279746A1

Abstract

A process is disclosed for selectively leaching lead sulphide from a lead sulphide bearing material (101). The process comprises a leaching stage (100) in which 5 the lead sulphide bearing material is mixed with an aqueous halide solution (303). The aqueous halide solution has an acidity, concentration and redox potential whereby the lead sulphide is selectively leached to produce a solubilised lead halide (103) and whereby hydrogen sulphide gas (104) is produced. The process also comprises passing the solution comprising the solubilised lead halide (103) to a precipitation stage (200) in 10 which the hydrogen sulphide gas (104) is recombined with the solution, whereby lead sulphide is precipitated (202). Optionally, the process further comprises passing the solution from the precipitation stage (200) to a regeneration stage (300) in which the acidity in the aqueous halide solution is regenerated such that the solution is able to be recycled (303) to the leaching stage (100). 15 ca -0 M 2 2 E E E U)U) U) 1Cl LLl E 04 C/)&) 0) -J a U) 0) 0C/) l 0 a, M o

Description

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Beneficiation of Lead Sulphide Bearing Material

Technical Field A process is disclosed for the beneficiation of galena, and other lead sulphide bearing materials and minerals, including ores, concentrates, tailings, and other such materials or residues. Such materials can be in the form of flotation products, filter cakes, mined ore, tailings, etc. The process can be applied to selectively separate the lead component from other metal components that can be present in the material, such as zinc, copper, nickel, cobalt, silver, gold, etc.

Background Art A known method for processing mineral ores is to float the target minerals using froth flotation. When the ores contain mixtures of lead, copper, zinc, silver, gold, bismuth, antimony, etc then differential flotation techniques tend to be used. However, in some situations, it is not possible to adequately separate the target metals from one another and produce individual metal concentrates. Typical mixtures of metals which can be hard to differentially float include lead with copper, lead with copper and zinc, lead with gold, and lead with zinc. Downstream processing of concentrates is designed to treat the main metal component, with additional metals adding complexity, and thus incurring penalties or limited payments to the concentrate seller. Examples include zinc smelters, where secondary circuits are required for lead processing, or copper smelters where secondary circuits are required for silver and gold processing. Lead sulphide is often processed using pyrometallurgical smelting techniques, with the secondary metals reporting to the slag component. The Imperial Smelting Process was designed to handle mixed lead-zinc sulphides, however, the zinc produced contains more trace metal contaminants than electrolytic zinc from a zinc roast-leach electrowinning refinery. Some hydrometallurgical processes have also been proposed for treating lead containing mixed-metal sulphide bearing materials. These include sulphuric acid pressure leaching, and halide systems, oxalic acid (US Patent 2839390) and, more recently, citrate acid or hydrogen peroxide systems (Zarate et al; Canadian Journal of Metallurgical Quarterly, Vol. 54, 2015 "Selective leaching of lead from a lead-silver zinc concentrate with hydrogen peroxide in citrate solutions"). Of the halide systems, some propose leaching the lead sulphide with oxidants and converting the lead into lead chloride crystals, and then treating the lead chloride crystals to recover lead metal or similar products. Alternatively, the lead can be leached and then directly recovered from the leachate by electrolysis. Examples of these systems include US patents 1435891, 1441063, 3929597, 4063933, 4082629 and CN201712488. US Patent 4024218 describes a process for converting the galena portion of a mixed sulphide concentrate into lead sulphate. The remaining copper sulphides can then be separated by flotation from the lead sulphate. However, the lead sulphate is undesirable for further treatment in a smelter, as the sulphate component is difficult to process. US Patent 7736606 describes a process for leaching of a value metal (nickel, copper, zinc, cobalt and Platinum Group Metals) from a base metal sulphide ore (pyrrhotite, pentlandite, chalcopyrite, arseno- and other pyrites, and sphalerite). The leaching process uses a combination of hydrochloric acid and an oxidant (alkali metal peroxide, alkali metal perchlorate, ammonium perchlorate, magnesium perchlorate, alkali metal chlorate, alkaline earth metal perchlorate, chlorine, alkali metal hypochlorite, hydrogen peroxide and peroxysulfuric acid). US 7736606 is silent on the leaching of galena or the like, nor does it mention lead recovery. In contrast, US 7736606 teaches that it is essential to use the oxidant as well as oxidising leach conditions to maintain the desired oxidising redox potential to thereby leach the target value metal from the base metal sulphide ore. The above references to the background art do not constitute an admission that the art forms a part of the common general knowledge of a person of ordinary skill in the art. The above references are also not intended to limit the application of the process as disclosed herein.

Summary of the Disclosure A process is disclosed for selectively leaching a lead sulphide bearing material (for example galena). The process can, for example, produce a high grade lead sulphide precipitate. The process as disclosed herein comprises (a) a leaching stage in which the lead sulphide bearing material is mixed with an aqueous halide solution that has an acidity, concentration and redox potential whereby the lead sulphide is selectively leached to produce a solubilised lead halide and whereby hydrogen sulphide gas is also produced. In contrast to the oxidising conditions taught in US 7736606, in the process as disclosed herein it has surprisingly been discovered that an oxidant is not required to selectively leach lead in a lead sulphide bearing material (such as galena). In fact, in the process as disclosed herein, it is desirable not to use an oxidant, such as those taught in US 7736606, because otherwise much or all of the other metal components present in the material (such as zinc, copper, nickel, cobalt, silver, gold, etc) will also be leached. Typically, in the leaching stage (a) the aqueous halide solution has sufficient acidity and sufficient ability to complex with the lead in the lead sulphide such that solubilised lead halide (e.g. lead chloride) is produced. This ability to complex can be achieved by optimising the solution composition (i.e. the halide composition is optimised so as to enhance lead complexation and solubility). In one embodiment, in leaching stage (a), the leach residual material (i.e. the material from which the lead sulphide has been selectively leached) may be separated from the solution (e.g. by filtering). The process as disclosed herein further comprises (b) passing the solution from (a) to a precipitation stage in which the hydrogen sulphide gas produced in (a) is recombined with the solution whereby lead sulphide is precipitated. In one embodiment, in precipitation stage (b), hydrogen sulphide from the leaching stage (a) may be supplemented with a complementary sulphidising reagent. This supplementing may be employed, for example, where insufficient hydrogen sulphide is generated in the preceding leaching stage (a). This in turn may arise when the lead sulphide bearing feed material comprises some non-sulphide lead components, such as lead oxide, lead sulphate, etc. The complementary sulphidisation reagent may be calcium sulphide, or may also be selected from calcium polysulphide, hydrogen sulphide, sodium sulphide, and/or sodium hydrogen sulphide. The process as disclosed herein brings together reagent production (i.e. hydrogen sulphide gas is produced in the leaching stage (a)), and reagent consumption (i.e. the hydrogen sulphide gas produced in (a) is reused (consumed) in the precipitation stage (b)). As a result, process economics can be considerably improved in comparison to known processes in which sulphidisation reagents (e.g. sodium sulphide) are purchased, or where biogenically derived hydrogen sulphide needs to be generated on site. In contrast to US 4024218, in the process as disclosed herein a high grade lead sulphide may be produced, and this may then easily be recovered in a known manner (such as by smelting and refining in primary lead smelters). Thus, the process has further advantages over the prior art where the lead product is typically lead chloride or lead sulphate. The process as disclosed herein may also be used to beneficiate ores and concentrates that contain mixtures of lead sulphide with minerals containing copper, zinc, silver, gold, cobalt, nickel, bismuth, antimony, etc. The process as disclosed herein, by selectively removing the lead sulphide component from the feed material, can also produce a residual material containing higher concentrations of the unleached metals (for example copper), being a residual material that may more easily be treated by known processes than when it contained the lead sulphide component. In one embodiment, the process as disclosed herein may further comprise a regeneration stage (c) to which the solution from (b) can be passed. In the regeneration stage (c) the acidity of the aqueous halide solution can be regenerated. In addition, the regeneration stage (c) may be operated such that an optimal solution redox potential can be achieved. The resultant aqueous halide solution may then be recycled to the leaching stage (a). In one embodiment, the acidity in the aqueous halide solution may be regenerated by the addition of an acid to the regeneration stage (c). For example, the acid that is added to regeneration stage (c) may comprise sulphuric acid. When, for example, the solution halide comprises chloride and/or bromide then hydrochloric acid and/or hydrobromic acid can be regenerated.

Thus, when the process comprises regeneration stage (c), it can be operated in a more efficient closed loop configuration. In this regard, the acidic aqueous halide solution that leaches the primary lead sulphide in leaching stage (a) can be subjected to sulphidisation in the precipitation stage (b) and can then be regenerated in stage (c) for recycle to and reuse in the leaching stage (a). In one embodiment, a neutralising agent, such as a metal alkali, may be added to the precipitation stage (b). This neutralising agent can cause the lead sulphide to be precipitated. The neutralising agent may, for example, comprise limestone, lime, sodium carbonate, sodium hydroxide, magnesium carbonate, magnesium hydroxide, magnesium oxide, etc. In one embodiment, the acid that is added to the regeneration stage (c) (e.g. sulphuric acid) may be selected such that its anion forms a precipitate with the cation of the neutralising agent. This precipitate may then be separated (e.g. filtered) from the solution in regeneration stage (c). For example, when the neutralising agent is e.g. limestone or lime, and the acid is sulphuric acid, the precipitate produced in regeneration stage (c) can comprise calcium sulphate. In one embodiment, the temperature of solution in the acid generation stage (c) can be controlled to determine the mineral form of the precipitated calcium sulphate, for example gypsum, basanite or anhydrite. As set forth below, the solution temperature in acid generation stage (c) may be in the range 20-115 °C, optimally around 80 °C. In one embodiment, the redox potential (Eh) in leaching stage (a) may be controlled to be in the range of -1000 to 300 mV (ref. Ag/AgCl). More particularly, the Eh may be in the optimal range of -50 to 50 mV (ref. Ag/AgCl). More specifically, the Eh may be less than around 30 mV (ref. Ag/AgCl). Optimally, the Eh may be controlled to be around 0 mV (ref. Ag/AgCl). Such a range and values of Eh can enable the lead sulphide to be selectively leached (i.e. in preference to other metal sulphides). At the same time, control of solution Eh can ensure that the sulphur component of the lead sulphide is converted into hydrogen sulphide and is not oxidised to elemental sulphur or sulphate. In one embodiment, the solution pH in leaching stage (a) may be less than 7. Optimally, the solution pH in leaching stage (a) may be controlled to be in the range of

0-1. Again, such a range and values of pH can enable the lead sulphide to be selectively leached whilst, at the same time, enabling the hydrogen sulphide gas to be produced. In one embodiment, the solution temperature in leaching stage (a) may be in the range 25-115 °C. Optimally, the solution temperature in leaching stage (a) may be around 80 °C. In one embodiment, leaching stage (a) may be operated at atmospheric pressure. In one embodiment, the residence time of the lead sulphide bearing material in leaching stage (a) may range from 0.1-24 hours. Optimally the residence time may be around 1-2 hours. In one embodiment, the solution pH in precipitation stage (b) may be less than 7. For example, the solution pH in precipitation stage (b) may be controlled (e.g. by the addition of the neutralising agent) to be in the range of 1-4. Optimally, the solution pH may be controlled to be around 2. Such a range and values of pH can favour the precipitation of stable lead sulphide. In one embodiment, the solution temperature in precipitation stage (b) may be in the range 20-115 °C. Optimally, the solution temperature in precipitation stage (b) may be around 70 °C. In one embodiment, precipitation stage (b) may be operated at atmospheric pressure. In one embodiment, the residence time of the solution in precipitation stage (b) may range from 0.5-24 hours. Optimally, the residence time may be around 1-2 hours. In one embodiment, the solution pH in acid generation stage (c) may be less than 7. Optimally the solution pH in acid generation stage (c) may be controlled to be in the range of 0-1. In one embodiment, the solution temperature in acid generation stage (c) may be in the range 20-115 °C. Optimally the solution temperature may be around 80 °C. In one embodiment, the acid generation stage (c) may be operated at atmospheric pressure. In one embodiment, the residence time of the solution in acid generation stage (c) may range from 0.5-24 hours. Optimally, the residence time may be around 1-2 hours.

In an alternative embodiment, acid generation stage (c) may be conducted in a pressure vessel (e.g. autoclave). The solution temperature in the pressure vessel may be in the range 115-200 °C. The pressure may be in the range of1-10 ATM. In one embodiment, the aqueous halide solution concentration may be in the range 1-10 moles per litre of solution. Optimally, the aqueous halide solution concentration may be controlled to be around 5 moles per litre. In one embodiment, the aqueous halide solution may comprise a metal halide solution. The metal halide may be one or more of: NaCl, NaBr, CaC 2, and CaBr 2

. However, it should be understood that the use of other alkali metal halides is possible, such as e.g. MgCl 2 .

In a further embodiment of the process, the selectively leaching of lead sulphide may be accompanied by different leaching, flotation or other separation stages for other target metals, such as may be present in the leach residue. Examples may include oxidative leaching of copper sulphide minerals, gold leaching using cyanide, froth flotation of the lead leach residue, and other known manners for treating metal bearing materials. Accompanying stages for target metals other than lead, are beyond the scope of the presently disclosed process, and constitute examples where the disclosed process can be operated within a multi-metal refinery or processing plant. In one embodiment, each of the leaching stage (a), precipitation stage (b) and regeneration stage (c) may be provided as circuits. Further, these circuits may be integrated. In addition, the leaching stage (a), precipitation stage (b) and regeneration stage (c) may each comprise multiple reactor stages. Employing multiple reaction stages can allow for better control of each of the stages (a), (b) and (c), generally resulting in improved yields, and better targeting of specific impurities or to-be recovered metals. The multiple reaction stages may each be operated in a co-current configuration. A co-current configuration can allow for better integration of the flow circuits between the leaching stage (a), precipitation stage (b) and regeneration stage (c), with minimal or simple solid/liquid separation equipment required. However, in some applications of the process, a counter-current configuration may be adopted for the leach circuit. This configuration may be required where the specific feed materials are complex and the counter-current configuration can assist the selective leaching of component minerals and chemical compounds. In a variation of the process, in leaching stage (a), other metal sulphides may be simultaneously leached with the lead, such as zinc and silver. The extent of leaching of these metals can be a function of mineralogy, temperature, and available acid. These metals can be recovered together with lead in the primary precipitation stage (b), or in additional metal recovery stages.

Brief Description of the Drawings Notwithstanding any other forms which may fall within the scope of the process as defined in the Summary, specific embodiments will now be described, by way of example only, with reference to the accompanying drawing in which: Figure 1 shows a block diagram for an embodiment of the process comprising a number of circuits that are integrated to process mixed metal sulphide concentrates and product high grade precipitated lead sulphide.

Detailed Description of Specific Embodiments In the following detailed description, reference is made to the accompanying drawing which forms a part of the detailed description. The illustrative embodiments described in the detailed description, depicted in the drawing and defined in the claims, are not intended to be limiting. Other embodiments may be utilised and other changes may be made without departing from the spirit or scope of the subject matter presented. It will be readily understood that the aspects of the present disclosure, as generally described herein and illustrated in the drawing can be arranged, substituted, combined, separated and designed in a wide variety of different configurations, all of which are contemplated in this disclosure. Flowsheet Description Figure 1 shows a process flowsheet in block diagrammatic form. The flowsheet illustrates one embodiment for the selective leaching of lead sulphide from an ore, concentrate, tailings or other lead sulphide bearing material. The flowsheet comprises three main integrated circuits, a leach circuit 100, followed by a lead sulphide precipitation (or sulphidisation) circuit 200, and then an acid generation (or regeneration) circuit 300.

Additional circuits for recovery of other metals can be included, such as further precipitation stages, solvent extraction, and/or ion-exchange resins, as may be the case for recovering leached metals which were leached either simultaneously or in separate stages to the lead sulphide leach. The flowsheet depicts a closed-loop process whereby the leaching circuit 100, precipitation circuit 200 and regeneration circuit 300 (i.e. stages) are integrated to bring together reagent production, regeneration and consumption. This can improve process flow characteristics and process economics, including the deployment of capital. Firstly, the lead sulphide bearing material is leached in the leach circuit 100 using acid in a salt brine, with the metal typically being calcium (although it could comprise magnesium or sodium). A typical acid employed in the leach circuit 100 is a hydrohalic acid such as hydrochloric and/or hydrobromic acid (i.e. such as is regenerated in circuit 300). In the process as depicted, the concentration of the halide solution can be in the range of 1-10 moles per litre of solution, optimally around 5 moles per litre. In leach circuit 100 the lead-free leach residues are recovered by filtration and washing to produce leach residue stream 102. Hydrogen sulphide gas (H2S(g)) is able to generated in the leach tanks, because in leach circuit 100 no air or other oxidant is injected into the process solution, hence the solution redox conditions are what can be referred to as "reducing". Hydrogen sulphide gas also has a limited solubility, and is therefore able to be captured from the headspace of the leach reactors to be directed to the precipitation reactors as stream 104. Secondly, the solution comprising solubilised lead halide (e.g. lead chloride and/or lead bromide) is passed as stream 103 to the precipitation circuit 200. In circuit 200 the precipitation of lead sulphide occurs as the hydrogen sulphide is re-injected into the lead-rich leach solution at an elevated pH. The elevated pH is controlled such as by the addition of a neutralising agent stream 201 (e.g. limestone or lime) to be in the range of 1-4, more typically around pH 2. In circuit 200 the sulphide precipitate is recovered such as by filtration to produce a high-grade lead sulphide ("artificial galena") stream 202. The filtered lead sulphide is washed, and packaged for sale.

Thirdly, the now lead-free salt solution 203 is passed to the regeneration circuit 300. In circuit 300 the acidity and redox potential of the halide solution is regenerated, such as by the addition of sulphuric acid stream 301. Where the neutralising agent stream 201 comprises limestone or lime, the addition of sulphuric acid produces a calcium sulphate precipitate, which is removed as stream 302, which is filtered and washed for sale/disposal. The addition of sulphuric acid also regenerates the hydrohalic acid (e.g. hydrochloric acid (HCl) and/or hydrobromic acid (HBr)). The HCl (and/or HBr) solution is returned (recycled) to the leaching circuit 100 as stream 303. The amount of sulphuric acid added is controlled to achieve an optimal redox potential Eh in stream 303. The typical chemistry (reactions - Rn) of the three circuits (100, 200, 300), assuming that the carrier brine contains calcium chloride, can be summarised as follows:

Leaching PbS()s+ 2HCl(aq)- PbCl 2 (aq)+ H2S(g) Rn 1 Sulphidisation PbCl2(aq)+ H2S(g) PbSts>+ 2HCl(aq) Rn 2 Acid Regeneration H2SO4(aq)+ CaCl 2 .nH 2 0(aq)- CaSO 4 .nH 20(s) + 2HCl(aq) Rn 3

Each of the circuits 100, 200 and 300 can comprise one or more recycle streams to allow for control of solids residence time in the solution as well as to improve yield/recovery. Each recycle stream can be from a given reactor stage to a previous reactor stage; a so-called "internal" recycle (for example the slurry from one reactor is recycled back to a previous reactor). Alternatively or additionally, each recycle stream can be from a separation stage (see below - for example from a thickener slurry underflow) to a given reactor stage; a so-called "external" recycle. Leach Circuit 100 (in detail)

The leach circuit 100 usually comprises a series of reactors followed by a separation stage to collect the leach solution and leach residue. The number of reactors, thickeners, filters, recycle streams, and other process equipment can vary depending on the complexity of handling of the solid material being leached.

The feed material (stream 101) is mixed with the hydrohalic acid (stream 303) recycled from the acid generation circuit 300. The slurry density range is typically from 0.5-50% w/w, and is often adjusted to maximise lead halide solubility. The oxidation reduction potential is typically maintained at < 300 mV (versus Ag/AgCl) to prevent oxidative leaching of the sulphide minerals, and to ensure that H2S(g) is released from the slurry. More specifically, the oxidation-reduction potential is maintained in an optimal range of -50 to 50 mV (ref. Ag/AgCl), typically less than around 30 mV (ref. Ag/AgCl), and optimally around 0 mV (ref. Ag/AgCl). At these levels the oxidation reduction potential is, in effect, "reducing" such that typically just the lead sulphide portion of the feed material 101 is leached (i.e. selectively) into solution. If the potential rises above 300 mV, then the sulphur component in the sulphide minerals will tend to oxidise to elemental sulphur or sulphate anions, and other (e.g. unwanted) metal sulphides may be leached. Additional, subsequent reactors can employ oxidative leaching conditions to target other metals once the lead sulphide has been leached (e.g. in a first or early stages of leach circuit 100). For example, the subsequent reactors can employ oxidative conditions for e.g. copper leaching - i.e. after lead sulphide has already been selectively leached. In leach circuit 100, leaching is carried out at a temperature in the range of 20 115°C, optimally at around 80°C, and for a residence time of 0.1-24 hours under atmospheric pressure. Often the lead sulphide leaches rapidly, and a residence time of < 2 hours (i.e. around 1-2 hours) can be sufficient. Sulphidisation Circuit 200 (in detail) The sulphidisation circuit 200 usually comprises a series of reactors followed by a separation stage to collect the leach solution (stream 103) and leach residue (stream 102). The number of reactors, thickeners, filters, recycle streams, or other process equipment can vary as determined by the production throughput. In sulphidisation circuit 200 the leach solution 103 mixes with the hydrogen sulphide gas stream 104 that is recovered from the leach circuit 100. Typically, the gas stream 104 also contains water vapour and a carrier gas (such as nitrogen, air, etc) to dilute the hydrogen sulphide for safety purposes.

Precipitation in the sulphidisation circuit 200 is carried out at a solution pH of less than 7, and typically in the range of 1-4 (optimally around pH 2) being a range/value that favours production of stable target metal sulphide precipitates. The temperature can be between 20-115 °C, and optimally around 70 °C. The solution residence time can range from 0.5-24 hours at atmospheric pressure, and is typically around 1-2 hours. The pH is adjusted through the addition of a metal alkali neutralising agent (stream 201), or a mixture thereof, selected from one or more of the following: calcium carbonate (e.g. limestone), calcium hydroxide and calcium oxide (e.g. lime), sodium carbonate, sodium hydroxide, magnesium carbonate, magnesium hydroxide, and magnesium oxide. The solution pH is controlled to provide the ability to selectivity precipitate dissolved lead from the solution, including the situation where the solution contains other metals, such as zinc, copper, iron. Optimal conditions can be adjusted as appropriate be persons skilled in the art. Acid Generation Circuit 300 (in detail) The acid generation circuit 300 usually comprises a series of reactors followed by a separation stage to collect the regenerated acid solution and residual precipitated material. The number of reactors, thickeners, filters, recycle streams, or other process equipment can vary as required to recover any precipitated sulphates. In circuit 300 the lead-free solution stream 203 from sulphidisation circuit 200 is mixed with sulphuric acid stream 301. The reaction produces a metal sulphate (for example CaSO 4 or Na 2 SO 4 ) and a hydrohalic acid (usually hydrochloric and/or hydrobromic acid). Where the process is operated with calcium chloride as the basis of the brine solution, this will most easily remove the sulphate added into the solution by sulphuric acid addition, in the form of precipitated calcium sulphate. This is because the solubility of calcium sulphate is much lower than sodium or magnesium sulphate. However, if there is no calcium present in the brine solution, then the corresponding sulphate salt (e.g. sodium or magnesium sulphate) can be recovered in a known manner (for example, by evaporation, crystallisation, ion-exchange, etc). The acid generation circuit 300 is usually operated at atmospheric pressure and at a temperature in the range of 20-115 °C (optimally around 80 C), with the higher temperatures resulting in less waters of crystallisation deporting with the calcium sulphate. That is, at low temperature the mineral form is mainly gypsum (CaSO 4 .2H 20), with bassanite (CaSO 4 .0.5H 2 0) formed at intermediate temperatures, and anhydrite (CaSO4 ) at elevated temperatures. The solution pH would be <1 (optimally in the range of 0-1) on account of the hydrohalic acid concentration. Solution residence time in circuit 300 typically ranges from 0.5-24 hours, optimally around 1-2 hours. To increase reaction times, and to produce anhydrite (CaSO4), at least part of circuit 300 can be conducted in a pressure vessel such as an autoclave, operating at a pressure in the range of1-10 ATM and a temperature in the range 115-200 °C. The precipitated calcium sulphate is separated as stream 302, and the regenerated hydrohalic acid brine is returned (recycled) to the leach circuit as stream 303. Solids-Liquid Separation Appropriate flocculants and coagulants can be added to the slurries to improve the efficiency of the solid-liquid separation stages. Typically each separation stage comprises a thickener and a filter, but alternatives can be a counter-current decantation stage, a single stage filter, or similar equipment. The thickening stage can make use of high rate thickeners, low rate thickeners, clarifiers and similar devices for solid-liquid separation. The filtration stage can make use of pressure filters, pan filters, belt filters, press filters, centrifuge filters and similar devices for solid-liquid separation. The acid regeneration circuit 300 may not require a thickening stage. Typically, each slurry is first sent to a thickener; with the resulting underflow slurry then forwarded to a filter for recovery of solids. The overflow can comprise process solution, or may be further filtered. Washing of the solids during recovery is employed to minimise any losses of process solutions and salts from the circuit. Fresh water is required for washing, and this is evaporated in each of the process reactors in the leach, sulphidisation and acid generation circuits. The resulting water vapour is discharged through the off-gas scrubber system or condensed and recycled as fresh wash waters. Filtered and washed solids can, in the case of: Leaching circuit 100 - comprise the leach residue 102 that contains unleached metals and constitutes a beneficiated product that contains minimal lead sulphide (for example a substantially lead-free copper sulphide concentrate);

Sulphidisation (precipitation) circuit 200 - comprise the precipitated lead sulphide, which can be further processed in a known manner for recovery of the lead component (for example smelting); Acid regeneration circuit 300 - comprise the metal sulphate precipitate (e.g. calcium sulphate) representing a saleable or recyclable product or which can be disposed of. Off-gas handling and scrubbing Off-gases are transferred from the various process reactors. The leach circuit off-gas contains the hydrogen sulphide gas which is transferred for consumption in the precipitation circuit. Unused hydrogen sulphide gas is then transferred to the scrubber unit, where it is chemically treated in a known manner to prevent any emission to the surrounding air environment.

Examples Non-limiting Examples of various stages (circuits) of the process for selectively leaching lead sulphide, and recovering precipitated lead sulphide from sulphide concentrates will now be described. Example 1 - Selective lead sulphidisation using hydrogen sulphide gas generated from a simultaneously operating leach circuit Two different lead sulphide concentrates containing lead, copper, and zinc were leached in sequence to produce 34 L of leach solution for sulphidisation tests. The batch leaches were conducted at 80°C and in each case a total of 3.6 kg of feed concentrate (1.8 kg for each feed source) was processed. A total of 12 leaches were completed, with the resulting slurries filtered and the solutions forwarded to the sulphidisation circuit. The solution used for each leach comprised 177 g/L NaCl, 110 g/L CaCl2, and 8.8 g/L HCl. The sulphidisation circuit was operated at 400 C, with the incoming solution from one leach combined with the leach off-gas (containing hydrogen sulphide) from the following batch leach. In this manner, the sulphidisation circuit was operated simultaneously with the leach circuit, using generated hydrogen sulphide from the leach circuit off-gas for precipitation. The sulphidisation operations were conducted in parts, to illustrate the sulphidisation stage using different neutralising agents. The test data is summarised in Table 1.

Table 1: Leach Extractions from Concentrates C and D, and precipitated lead sulphide grades

Weight Metal Cu(%) Pb (%) Zn(%) Fe (%) Wkg (kg) Concentrate C Feed Grade 0.48 21.70 9.59 8.63 1.800 Leach Residue 0.67 2.32 8.74 11.30 1.176 Grade Leach Extraction 9% 93% 41% 15% Concentrate D Feed Grade 12.30 23.70 7.67 16.20 1.800 Leach Residue 18.18 0.74 7.67 23.00 1.165 Grade Leach Extraction 5% 98% 35% 8% Volume Cu (g/L) Pb (g/L) Zn (g/L) Fe (g/L) (L) Solution After <0.01 11.1 2.1 0.3 34 Leaching

Neutralising Agent Cu (%) Pb (%) Zn (%) Fe (%) Ca (%) Sulphidisation CaCO 3() 0.08 83.6 0.29 0.52 0.01 A Sulphidisation NaOH(aq) 0.01 83.4 0.02 0.04 0.01 B

The lead sulphide in the feed concentrates was selectively leached to an extent of 93-98%, in comparison to the copper (5-9%) and iron (8-15%). Approximately 35 41% of the zinc was also leached. These leaches produced a combined 34 L of leach solution containing 11.1 g/L lead, 2.1 g/L zinc, 0.3 g/L iron, <0.01 g/L copper. The sulphide precipitation was conducted at a pH of 2-3 in a batch reactor, with variations in the selection of neutralising agents. The first test was with the addition of calcium carbonate (limestone), the second test was with the addition of only sodium hydroxide. The grade of the precipitated lead sulphide was similar in both tests at 83.4 83.6%. However, the contaminant zinc, iron, and copper were slightly higher in the test with calcium carbonate than the test with sodium hydroxide. This difference is due to greater utilisation of the liquid reagent (sodium hydroxide) compared to the solid reagent (calcium carbonate). The results confirmed that the sulphidisation stage using the off-gas from the leach circuit stage, could be operated to selectively precipitate high grade lead sulphide from the leach solutions containing lead and zinc. Example 2 - Generation of hydrohalic acid for leaching using spent solutions after sulphidisation in Example 1 A 4 L portion of the spent sulphidisation solutions was collected after Example 1, and treated in an acid generation circuit. The solution was heated to 77°C, and 98 g of sulphuric acid was added. The reactor was mixed for 1 hour, and the resulting slurry was filtered to separate the acidic solutions and the precipitated calcium sulphate. The test data is summarised in Table 2.

Table 2: Treatment of spent sulphidisation solution with sulphuric acid

Zn Fe Ca HCl (g/L) Volume Metal Cu Pb (g/L) (g/L) (g/L) (g/L) (g/L) (L) Feed<.1(H 40 <0.01 2.78 1.95 0.55 30.22 <0.1(pH 4.00 Solution 3)

Exit Solution <0.01 2.95 2.06 0.59 22.60 19.4 3.75

Cu (%) Pb (%) Zn (%) Fe (%) Ca (%) Weight (g) Precipitate <0.01 <0.01 <0.01 <0.01 25.39 139

The precipitate was washed and dried at 50 °C, and the mineralogy identified by X-ray diffraction. The precipitate was shown to be a mixture of 82% basanite (CaSO4 .0.5H20) and 18% gypsum (CaSO 4.2H 20). There was negligible precipitation of other metals.

The final solution contained 19.4 g/L of HCl, as the sulphate cations had been precipitated with the calcium. This solution was then forwarded to a leach circuit, completing the integrated stages of the process.

Whilst a number of specific process embodiments have been described, it should be appreciated that the process may be embodied in other forms. In the claims which follow, and in the preceding description, except where the context requires otherwise due to express language or necessary implication, the word "comprise" and variations such as "comprises" or "comprising" are used in an inclusive sense, i.e. to specify the presence of the stated features but not to preclude the presence or addition of further features in various embodiments of the process as disclosed herein.

Claims

Claims 1. A process for selectively leaching lead sulphide from a lead sulphide bearing material, the process comprising: (a) a leaching stage in which the lead sulphide bearing material is mixed with an aqueous halide solution that has an acidity, concentration and redox potential whereby the lead sulphide is selectively leached to produce a solubilised lead halide and whereby hydrogen sulphide gas is produced; (b) passing the solution from (a) to a precipitation stage in which the hydrogen sulphide gas produced in (a) is recombined with the solution whereby lead sulphide is precipitated.
2. A process as claimed in claim 1 wherein a neutralising agent, such as a metal alkali, is added to the precipitation stage (b) whereby the lead sulphide is precipitated.
3. A process as claimed in claim 1 or 2 further comprising: (c) passing the solution from (b) to a regeneration stage in which the acidity in the aqueous halide solution is regenerated such that the solution is able to be recycled to the leaching stage (a).
4. A process as claimed in claim 3, wherein the acidity in the aqueous halide solution is regenerated by the addition of an acid, such as sulphuric acid, to regeneration stage (c).
5. A process as claimed in claim 4, when dependent on claim 2, wherein the acid is selected such that its anion forms a precipitate with the cation of the neutralising agent, which precipitate is separated from the solution in regeneration stage (c).
6. A process as claimed in any one of the preceding claims, wherein the redox potential (Eh) in leaching stage (a) is controlled to be in the range of -1000 to 300 mV (ref. Ag/AgCl), optimally in the range -50 to 50 mV (ref. Ag/AgCl), to enable the lead sulphide to be selectively leached.
7. A process as claimed in claim 6, wherein the redox potential (Eh) in leaching stage (a) is controlled to be around 0 mV (ref. Ag/AgCl).
8. A process as claimed in any one of the preceding claims, wherein solution pH in leaching stage (a) is less than 7, optimally controlled to be in the range of 0-1.
9. A process as claimed in any one of the preceding claims, wherein solution temperature in leaching stage (a) is in the range 25-115 °C, optimally around 80 °C.
10. A process as claimed in any one of the preceding claims, wherein leaching stage (a) is operated at atmospheric pressure, and wherein the residence time of the lead sulphide bearing material in leaching stage (a) ranges from 0.1-24 hours, optimally around 1-2 hours.
11. A process as claimed in any one of the preceding claims, wherein solution pH in precipitation stage (b) is less than 7, optimally controlled to be in the range of 1-4 to favour the precipitation of lead sulphide.
12. A process as claimed in claim 11, wherein the solution pH in precipitation stage (b) is controlled to be around 2.
13. A process as claimed in any one of the preceding claims, wherein solution temperature in precipitation stage (b) is in the range 20-115 °C, and optimally around 70 °C.
14. A process as claimed in any one of the preceding claims, wherein precipitation stage (b) is operated at atmospheric pressure, and wherein the residence time of the solution in precipitation stage (b) ranges from 0.5-24 hours, optimally around 1-2 hours.
15. A process as claimed in claim 3, or any one claims 4 to 14 when dependent on claim 3, wherein solution pH in acid generation stage (c) is less than 7, optimally controlled to be in the range of 0-1.
16. A process as claimed in claim 3, or any one claims 4 to 15 when dependent on claim 3, wherein solution temperature in acid generation stage (c) is in the range 20 115 °C, optimally around 80 °C.
17. A process as claimed in claim 3, or any one claims 4 to 16 when dependent on claim 3, wherein acid generation stage (c) is operated at atmospheric pressure, and wherein the residence time of the solution in acid generation stage (c) ranges from 0.5 24 hours, optimally around 1-2 hours.
18. A process as claimed in claim 3, or any one claims 4 to 15 when dependent on claim 3, wherein acid generation stage (c) is conducted in a pressure vessel at a solution temperature in the range 115-200 °C and at elevated pressures in the range of 1-10 ATM.
19. A process as claimed in any one of the preceding claims, wherein the aqueous halide solution concentration is in the range 1-10 moles per litre of solution, optimally controlled to be around 5 moles per litre.
20. A process as claimed in any one of the preceding claims, wherein the aqueous halide solution comprises a metal halide solution such as one or more of: NaCl, NaBr, CaC1 2, and CaBr 2

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